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By T Purvis, Z Liu

Ernest Henry mine (EHM) is an underground sublevel caving mine located in north-west Queensland with a production rate of 6.4 Mt/a. The mine has sublevels spaced with a 25 m vertical interval. The production process on each level starts with the drilling and firing of a blind uphole rise to start the slot extraction. These rises are fired as a single shot and routinely need to achieve a design height of 22 m to 30 m depending on the local mine geometry. If the rise fails to perform as designed then the production phase of the level is delayed with significant rework required to remedy the situation.This paper describes the process that the EHM team has followed to improve the first time performance of each rise. The improvement process commenced with a review of results from previous rises and resulted in the identification of a number of improvement opportunities. A revised holistic approach was established for the process, which included involvement from operators and engineers. Through continued measurement and review of rise performance, inputs to the process were improved to a point where reliable results can now be expected. The successful performance of subsequent rises has been verified through a comprehensive quality assurance quality control (QA/QC) program. Before quality control processes were implemented, the success rate of blind uphole rises was only 11 per cent (one out of nine rises) at EHM. In comparison, the success rates have been vastly improved to 25 per cent (one out of four) and 75 per cent (three out of four) respectively following two stages of improvements. This paper concludes with the ‘current’ series of best practices used at EHM and recommendations.CITATION:Purvis, T and Liu, Z, 2017. Improving the performance of blind uphole rises at Ernest Henry Mine – a case study, in Proceedings 13th AusIMM Underground Operators’ Conference 2017, pp 335–340 (The Australasian Institute of Mining and Metallurgy: Melbourne).

Oct 16, 2017

By V. E. Ross

A computer program was developed that incorporates a novel analysis of the behaviour of the froth in full-scale cells to predict the performance of the froth over a wide range of operating conditions. The program has the distinct advantage that it can be easily implemented on a microcomputer, which allows rapid computation and processing of the results. The program utilizes experimental data obtained in flotation-plant and laboratory testwork to provide parameters for the froth model under an initial set of operating conditions. These parameters can then be used in conjunction with a model describing the transfer of material from the pulp to the froth to predict the performance of the cell when its dimensions and operating variables are changed during scale-up.

Jan 1, 1987

By R. T. Carter

THE phase behavior of a number of binary systems consisting of paraffin hydro-carbons has been determined in recent years. In addition, the composition of the coexisting phases of mixtures of crude oil and natural gas has been investigated.1 However, the latter study did not include information concerning the influence of the concentration of a component upon its distribution between two coexisting phases. Owing to the paucity of such information it was considered desirable to investigate in some detail the composition of the co-existing phases in the methane-propane-n-pentane system throughout the greater part of the two-phase region at a tempera-ture of 100°F. Information of this nature also permits the establishment of the critical behavior of this ternary system.

Jan 1, 1940

By Donelson A. Robertson

Includes a review of the state-of-the-art of explosives loading in the crushed stone and cement industries of South Florida and the development of a field test program for alternatives. Emphasis is on the loading of small diameter (2 1/4" or less) explosives. Test results ire evaluated for operational and economic feasibility. Details are described of an explosives loading method which involves placing cardboard tubes with external couplings through the hollow drill steel and then Loading explosives into the tube casings after the drill rig has moved off. This method eliminates the hazards peculiar to loading explosives Through the kelly bar. The alternative method can be performed by existing drill rigs with or without minor modification. With modification, improvement in yield per blasthole and/or primary crushing rate provides opportunity so offset some or all of the additional cost.

Jan 1, 1978

By A. D. Schwope, L. R. Jackson, K. F. Smith

Burghoff and Blank1 have pointed out that the creep properties of hard-drawn coppers are closely associated with their individual softening characteristics and have further shown that the creep resistance of various types of copper is improved, under certain circumstances, by cold stretching. In the design of certain types of electrical machinery, it is important to have the creep resistance as high as possible; furthermore, short-time creep strength is important since in many cases stresses are large only during warming-up and cooling-down periods. In this paper, data are presented showing the effect of cold work on the short time creep strengths of various types of commercial copper under several conditions. It is shown that trends from short-time creep tests are evident at longer times. Materials Used in the Investigation The coppers used in this investigation were commercial grade and of two basic types, oxygen free high conductivity (OFHC) and tough pitch. Additions of silver in amounts of 15 and 25 oz per ton were added to each of the two coppers. The analyses of these six coppers are given in Table 1. Originally, they started as wire bars 4 X 4 X 54 in. A. square rod was made by hot rolling the wire bars to 9/16-in. rod in 8 passes. This rod was then cold drawn to 0.445-in.-diam rod and annealed for 3 1/2 hr at 1000°F. It was then cold drawn to 0.325-in. rod and again annealed at 1000°F for 3 1/2 hr. Hard square rods were then made by cold rolling to 0.257 X 0.257-in. rod. Soft wire was made from the hard- drawn material by annealing at 650°F for 3 hr. The final rolling was performed using a set of 9-in. grooved rolls revolving at 30 rpm, which produced square wires. Reductions of approximately 5, 20, 30, and 40 pct were obtained using several passes in the same direction for the large reductions. The 5 pct reduction by drawing was accomplished using a steel die having 2 parallel bearing surfaces. The die angle was 7 1/2 degrees and the wires were lubricated with a paste of ceresin wax dissolved in carbon tetrachloride. The speed of drawing was 13 fpm. Since work was done on only two sides of the wire simultaneously, it was necessary to rotate it 90 degrees each pass to insure uniform deformation. The room-temperature tensile properties of the various tempers are given in Table 2. The grain size of the an- nealed 0.257-in. wires was uniform and averaged about 0.035-0.045 mm. Experimental Procedures for Short-time Creep Tests Standard creep-test frames (Fig 1) were used for the testing. Special fixtures were designed to apply axial compressive and tensile loads to the copper wires. The compression specimens were 1 in. long and the ends were ground flat and parallel. The creep in compression was indicated by the movement of reference marks s-ratched on intersliding platinum strips attached to the moving heads of the fixture. The movement was measured with an optical micrometer fitted with a filar eyepiece on which the smallest division was 0.00005 in. The tensile creep specirriens were 4 in. long which provided ample gripping surface for the jaw-type grips. The platinum strips were attached to a one-inch gauge length of the wire and the creep measured as previously described. The single-specimen furnaces were connected to individual power and control sources. The controlling equipment consisted of Foxboro controllers

Jan 1, 1950

By Daniel W. H. Su

A nearly 305-m- (1,000-ft-) wide sandstone channel was projected to run through the mid panel area of a 457-m- (1,500-ft-) wide longwall panel in a southwestern Pennsylvania coal mine. This sandstone channel was known to cause serious longwallface roof control problems in the mine?s B-panel area in 1998 and 1999; and most recently, it induced many large face cavities and caused a substantial production delay in the preceding panel. For the two B panels mined in 1998 and 1999, hydraulic fracturing of the massive sandstone was conducted to facilitate better caving, thus relieving cantilever compressive stress at the longwall face. Vastly improved longwall face condition was observed and four weeks of mining delay due to bad roof conditions were eliminated. The implementation of the hydraulic fracturing technique, results of underground observation and measurement, and results of numerical modeling were summarized in a paper presented in the 20th International Conference on Ground Control (Su et al., 2001). This article is a summary of the activities that took place during a four-month period (October 2011 through February 2012) at a Consol Energy longwall mine southwestern Pennsylvania to improve face conditions and safety of underground personnel during mining of the 457-m- (1,500-ft-) wide longwall panel. The elevated concern for safety was driven by adverse face conditions encountered while mining the preceding longwall panel. Many of the problems were attributed to a thick and massive sandstone channel. A joint task force consisting of coal operations engineering, Pennsylvania operations engineering and CNX Gas personnel was formed in August 2011, which recommended drilling four hydraulic fracturing holes along the center axis of the sandstone channel to reduce pressure and improve face conditions at the longwall face. To further provide detailed monitoring of the longwall face and to evaluate the effectiveness of the hydraulic fracturing program, the Longwall Visual Analysis (LVA) software (Hoyer, 2011) was installed in early December 2011 to track the face pressure and c

Jun 1, 2013

By Daniel Lachapelle, Nuri Hmidi, Turgut Yalcin

"Simulation can serve as a convenient tool in studying the behaviour and performance of grinding circuits by facilitating the evaluation of various scenarios and identification of optimum operating conditions. This technique was applied to the St Andrew Goldfields Ltd.’s Stock Mine Mill which is a 1300 tpd gold milling facility, offering custom milling services as well as processing the Company’s own ores. Lack of adequate instrumentation in the mill poses a challenge in attaining optimum performance especially with changing ore types and characteristics. The grinding process is monitored essentially by manual measurement and adjustment of pulp densities at various streams. A simulation program was designed to assist in the effective implementation of this mode of grinding control. Product sizes from both the primary and secondary grinding circuits were correlated empirically to relevant process variables using a database formed from the results of past sampling campaigns. The simulator was created in Microsoft Excel whose data analysis and optimization tools coupled with its Visual Basic Programming capabilities were found to provide an excellent environment for the development of a cost-effective, efficient and user-friendly simulation program.INTRODUCTIONGrinding is often a very costly unit process in milling operations, mainly due to the excessive energy consumption associated with it. For example, for a Canadian copper-zinc flotation mill, grinding was listed as being the most costly item taking up 30.4% of the total milling expenses, followed by the flotation cost at 2 1 .O% (Kidd Creek Mines Ltd., 1982). A survey on 45 copper flotation mills around the world showed the average grinding cost to be 34.5% of the total milling cost (Weiss, 1985).Overgrinding aggravates the cost issue, while insufficient grinding either causes loss of valuable minerals to tails or results in low-grade concentrates. Therefore, it is essential that grinding circuits be carefully controlled and monitored. To this end, a great deal of effort has been devoted over the years to the modelling and simulation of grinding systems in order to assist in the efficient running of these processes."

Jan 1, 2001

By Fred N. Kissell, Maurice Deul, T. D. Moore

Methane emission from two surface gob-degasification boreholes was measured by the Bureau of Mines. The investigation took place during mining of a 3,200-foot longwall panel in the Lower Kittanning coalbed in central Pennsylvania. The first hole was 500 feet from the start of mining, and the second hole was 2,200 feet. The first hole went on natural draft as soon as the longwall face passed it. This caused the methane flow in the return entries to drop 75 pct. In the 6 months required to complete the panel, the two holes emitted 69 MMcf of methane in varying percentages of concentration in air. In 3 years, the holes emitted a total of 150 MMcf of methane.

Jan 1, 1976

By L. B. Ticknor, M. B. Bever

An isothermal calorimeter suitable for measurements of heats of solution in liquid tin as solvent is described. Measurements of the heats of solution of gold, silver, copper, and a gold-silver alloy are reported. The heat of formation of this gold-silver alloy is also reported. HEATS of solution in liquids can be determined directly, but few such data have been published for metals as solvents. Knowledge of heats of solution is desirable because it contributes to the general understanding of energy relations in metallic solutions. Thermochemical measurements of this type also furnish answers to specific problems: For example, from the heats of solution of an alloy and of a mixture of its constituent elements in a given solvent, the heat of formation of the alloy may be found. Similarly, differences in the energy content of samples of the same metal in two different states, such as the annealed and cold worked states, may be determined. In the work reported here, gold, silver, and an alloy of gold and silver were dissolved in liquid tin at 240°C and gold, silver, and copper were dissolved in tin at 300 °C. The experiments yielded the heats of solution of these metals in tin. The heat of formation of the gold-silver alloy was also determined. The same calorimetric method is being used to measure the energy stored in metals during cold working and preliminary results have been published.' Equipment and Experimental Procedure The isothermal calorimeter used in this work, Fig. 1, consisted of a Dewar flask completely immersed in a constant-temperature salt bath. A long neck provided space for thermocouple leads and a stirrer shaft and permitted the injection of solute samples. The metal bath in the calorimeter flask was stirred at about 100 rpm by a small glass impeller driven by a constant-speed motor. The shaft of the stirrer passed through a Wilson vacuum seal.2 All measurements were made under vacuum in order to eliminate temperature disturbances resulting from gas convection and to prevent the oxidation of tin. The calorimeter flask was joined by an all-glass connecting line to a gas-control system which made possible the use of a protective atmosphere of deoxidized and dried hydrogen. At the operating temperature of the calorimeter tin oxide was not reduced by hydrogen, presumably because the reaction rate was too slow. Part of the closed system was a removable side tube through which the solute sample could be introduced and in which it was held at a temperature of 0°C prior to injection into the metal bath. The salt bath containing a mixture of sodium, lithium, and potassium nitrates was similar to that described by Beattie3 and was equipped with a stir-

Jan 1, 1953

By Frank Robertson Van Horn

In June, 1911, the writer spent a few days in studying the economic geology of the vicinity of Park City. During this rather hurried visit a number of specimens of ore were collected, which have since been examind. The results of these investigations show that bournonite (PbCu-SbSa) is present in the ores in greater or less quantity. This fact has never before been reported, as far as can be learned. The failure to recognize this mineral is due to its similarity to tetrahedrite, with which it has probably been confused during a long period of years. Jamesonite (Pb2sb2s6) has been reported from the region in a rather doubtful manner, but the writer has seen it in notable amounts from several mines, and can offer the first analysis of this mineral from the district. Calamine (HzZn2SiO5) was found at one mine and has never been ecognized and reported from the district. Geogbaphy and Geology Park City, Utah, is located at an elevation of about 7,000 ft. above sea level on the eastern slope of the Wasatch mountains. It is situated approximately 25 miles southeast of Salt Lake City. ' The geology and ore deposits have been described in detail by Bout well and Woolsey.' The mines visited are situated south-southeast from the town. The Silver King Coalition mine is located in Woodside gulch about 1 mile southeast of Park City, at an altitude of about 8,100 ft. The other mines visited were the Daly West and Quincy, which are in Empire canyon about 2 miles sontheast of the town, at elevations of 8,300 to 8,500 ft. respectively. The Daly Judge tunnel goes into the mountain from Empire canyon about 1 mile south of the city, at an altitude of 7,700ft.

Jan 1, 1915

By N. J. Alderman, D. J. Clowes, N. I. Heywood

The effects were studied of the presence of comminuted flint on the rheological characteristics of chalk slurry pumped through a 92-km long pipeline from Kensworth Quarry to Rugby Cement Works in the UK, and their impact on the pipe flow conditions. Flow curves were measured for 63different chalk slurry samples; these were described by the Bingham plastic model. In addition to observing trends in the yield stress parameter and plastic viscosity in the model with variations in both flint concentration and dispersant additive, use was also made of slurry viscosity values at a reference shear rate of 60 s-1for the pipe flow. It was found that chalk slurry viscosity decreases with increasing flint content up to 10% by weight for a given additive concentration, and increases with increasing additive concentration from 0.7 to 1.0%, for a given flint content. The pipeline flow regimes were identified using the critical value of the Bingham Reynolds number for laminar flow breakdown. For Bench 11 (quarry bottom) chalk slurries, it was observed that all sections of the pipeline operate in the turbulent flow regime for the nine different formulations and for the three pipe diameters used along the pipeline. However, this was not always the case for Bench 1(quarry top) slurries. It was therefore recommended that chalk slurries prepared from Bench 1 should not be used because of the strong likelihood that laminar flow conditions will occur owing to the presence of higher clay concentrations within this bench. This, in turn, will lead to an unsafe operation of the pipeline because of the potential for coarse particle settling. Likewise, blending of chalk from Bench 1 with the lower benches such as Bench 11 could be problematic and this requires further experimentation.

Jan 1, 2007

By Herbert A. Everest

Discussion of the paper of HERBERT A. EVEREST presented at the New York meeting, February, 1916, and printed in Bulletin No. 109, January, 1916, pp. 165 to 167. NEWELL G. ALFORD, Earlington, Ky. (communication to the Sec-retary*).-Mr. Everest's statement concerning the economies in a small coal mine has interested me very much. The majority of discussions on coal-mine economy omit the small operator and for this reason Mr. Everest's paper should lead to productive thought. I agree that close concentration of working places and constructive supervision of labor are two items of salvation to the small operator. It is also true that small mines often yield as much profit per capital invested as their larger neighbors. Mr. Everest also says that the small operator has so keen a regard for the inroad on his profits that even one extra day hand is an extravagance not tolerated, and this is rightly so. But, does not this earmark of the small operator in many cases offer a clue to his vital weakness? A credit balance is the acid test of success in any enterprise. With the small operator who owns practically all the capital stock this is particularly significant. He never takes his eyes off the market price and at the same time strains every nerve to cut down his payroll. In several cases I have noted that the resultant close supervision of labor was in fact too close. The "stitch in time" was not applied to the repair of this or that, with a decided increase to the cost of the work when it had to be done at the expense of cutting off the run when orders were on the books.

Jan 5, 1916

By W. M. Courtis

This district commences near Pigeon River, the northeastern boundary between Minnesota and Province of Ontario, and extends entirely around the north shore of Lake Superior, terminating for the present at the Bruce Mines on Lake Huron. The discoveries of mineral-bearing veins have been confined for the most part to the lake shore, or to the country opened up by the Dawson Red River road. As yet the country back from the lake has been but little explored, on account of its roughness, being intersected by steep, rocky trap bluffs, 1000 or more feet high, and alternating white cedar and tamarack swamps. The forests are exceedingly difficult for prospectors. In many places Titanic piles of rocks, protected by chevaux-de-frise of dead cedars, with their hooked, tough, sharp branches, and a covering of spongy moss, sometimes several feet thick, successively guard against the discovery of hidden treasure. The Hudson's Bay Company for nearly a century have maintained trading-posts at Fort William and other points, which probably opened the way to the first discoveries that were made at Prince's Bay, near the western limit of the district. At this point a large vein was worked by Colonel Prince, the owner, as early as 1846. The richest part of the vein seems to have been on Spar Island, but work was stopped on account of excess of water. Sir William Logan states that a mass of ore weighing several hundred pounds and carrying 3 per cent. of silver was taken out. On the mainland small quantities of native and sulphide of silver were found, together with iron and copper pyrites, zinc blende, and galena, in a large spar vein. On the main shore a drift 165 feet long and 90 feet in depth

Acclaimed as one of the most remarkable orebodies anywhere on account of its length and richness, the Oroya Shoot occurs on the eastern side of the Kalgoorlie gold mining district close to the contact of a fine-grained with a coarse-grained greenstone, and in association with sediments. In the same stratigraphic position on the synclinal flanks of a large drag fold there are a number of other flat ore pipes, smaller, but almost as rich in tenor. Although differing in control structure they are believed to be related genetically as all show a characteristic development of a bright-green ore known locally as "green leader". Folding and bedding structures have played an important part in formation of the ore shoots.INTRODUCTIONThe Oroya Shoot is the biggest individual bonanza of Kalgoorlie's Golden Mile. It has been stoped almost continuously over a length of 4,300 ft., and about 2,000,000 oz. of gold have been won from it. Naturally pillars are not left in the best sections of such an orebody and the detailed geology of stoped out portions can only be inferred from examination of adjacent workings.No mention is made in the literature of the actual makeup of the orebody although a few structures are shown on some of the old cross sections. A diagrammatic cross section, Fig. 1, adapted from Stillwell (1929) illustrates the form of the Oroya Shoot.Gustafson (1937) drew attention to the role of folding, previously unrecognized, in Kalgoorlie geology, and indicated that the Oroya Shoot is associated with a minor drag fold on the east limb of the Brownhill Syncline, itself part of a larger left hand drag on the Central Anticline. The latter in turn appears to be part of a regional right hand fold system as shown by Campbell (1953).

Jan 1, 1959

By G. L. F. Powell

g samples of Cu-20 wt pct Ag alloy have been mdercooled to a maximum of 197°C by melting under a slag of commercial soda-lime glass in a vitreous silica crucible. No grain refinement of the primary copper was observed in samples undercooled to the maximum of 197°C. When the samples contained a small amount of oxygen, the copper dendrites were partially recrystallized at undercoolings greater than 97°C. In previous papers'-3 reporting the grain structure of undercooled silver and copper, it was observed that grain refinement was dependent on both undercooling and oxygen content. Grain refinement occurred in undercooled silver when the degree of undercooling exceeded the range 153" to 175"C, while in Ag-0 alloys (0.12 wt pct) fine equiaxed grains were exhibited when undercooling was greater than 50°C. Similarly, copper samples undercooled as much as 208°C displayed fan-shaped growth from a single nucleation site, while the grain structure of Cu-O alloys (0.08 wt pct) was fine and equiaxed at undercoolings larger than 150°C. Thus the presence of oxygen greatly reduced the undercooling at which grain refinement occurred. It was also observed that the change in grain size resulted from recrystallization and was not due to an enhanced nucleation rate in the liquid-solid transformation. It is possible that the influence of oxygen on recrystallization is due primarily to its presence as a solute element. walker4,' reported that, although a grain size change did not occur in pure nickel until the undercooling exceeded 150°C, small grains were observed in samples of Ni-Cu alloy solidified at small and large degrees of undercooling. Jackson et al.6 suggested that the fine grained structure of the Ni-Cu alloy resulted from the melting off of dendrite arms during recalescence. This remelting process may occur in alloys as a result of segregation during freezing which causes a variation in liquidus temperature from point to point within a dendrite. It was therefore decided to undercool copper with a metallic alloying element to ascertain whether the presence of a metallic solute would have a similar effect to oxygen in inducing grain refinement. A Cu-Ag alloy was chosen, since both metals had been shown to behave similarly on undercooling. The alloy Cu-20 wt pct Ag was selected since the eutectic constituent outlines the initial growth form of the primary copper, so that the as-frozen grain structure is not obscured if subsequent recrystallization occurs. This paper describes the results of undercooling experiments carried out with Cu-20 pct Ag samples undercooled to a maximum of 197°C and the effect of oxygen content on the grain structure of the undercooled samples. EXPERIMENTAL Melting was carried out in a small cylindrical resistance furnace using "fine" silver granulate and oxygen-free high conductivity copper. The procedure adopted was to melt the required quantity of silver in air in a clean vitreous silica crucible for approximately 15 min, freeze, and add granulated commercial soda-lime glass to form a complete surface slag cover, after which the sample was melted and frozen several times to reduce the oxygen content. The glass slag cover was approximately 3 in. thick. Pieces of copper (=50 g) were added to the crucible until the required quantity to make 350-g samples of alloy had been charged. Each piece was added quickly to the crucible which was held at a temperature slightly above the melting point of silver. The piece was quickly pushed beneath the glass to minimize oxidation and any oxide coating usually decomposed before the piece had settled down into the silver. After the full quantity of copper had been added, the melt was stirred with a silica rod to hasten homogenization and a Pt/Pt 13 pct Rh thermocouple enclosed in a vitreous silica sheath inserted for temperature measurement. Heating and cooling curves were recorded on a potentiometric chart recorder fitted with a zero suppression unit. The milli-voltage range of the recorder was adjusted so that temperatures could be read to 1°C. Heating and cooling curves were taken every hour until three consecutive readings gave the same solidus-liquidus range, consistent with the solidus-liquidus range for this alloy composition by reference to Hansen and Anderko.7 Metallographic examination of samples frozen at this stage, failed to show any variation in composition from bottom to top of the ingot. Consequently, it was considered that the melt was homogeneous at this stage and undercooling experiments were then car-

Jan 1, 1970

By L. J. Friedman

Smelter operating and environmental requirements demand high on-stream time (high reliability) from the primary smelter sulfur retention device - the sulfuric ?acid plant. The key to reliable smelter sulfuric acid plant operation is the gas cooling and cleaning system. Problems in the gas cleaning system cause corrosion, plugging, equipment failure, etc - plant down time, in the contact section of the plant. The last fifteen years have seen major changes in gas cleaning system design _to the current standard of brick lined venturi scrubber, packed gas cooling tower with plate type liquor coolers and lead or plastic tube precipitators, The current? design offers improved' impurity scrubbing along with increased reliability. in addition, a number of alternate gas cleaning systems have been offered Chemetica LAB System; Monsanto DuPont D:Jna-Wave Syatem1 Lurgi venturi scrubber with lead star gas cooler. This paper will review the design and operating advantages, disadvantages, and risk of the standard gas cleaning system and compare it to the alternate systems. _

Jan 1, 1992

By Amir Karami, Derek Kinakin, Catherine Schmid

The Mathews Stability Graph Method (Mathews et al., 1981) has been used by geotechnical engineers and mining engineers for over 40 years to estimate the performance of longhole open stopes in underground mines. The method is widely used and accepted by practitioners because it is relatively simple to use and can be modified to account for observed performance at a given mining operation. While the simplicity of the method increases its accessibility and utilization, it does not intrinsically account for variability in the various inputs, including rockmass quality, structure, stress, or stope orientation. This paper presents a modified approach for the Stability Graph Method, whereby Python scripts are used to process available geotechnical datasets and combine them with actual or potential distributions of Q parameters, induced stress, structure, or stope orientation. The resulting Stability Numbers are then plotted as a contoured distribution on the Stability Graph.  This allows the practitioner to quickly incorporate large datasets into the analyses, while also developing an understanding of how variability in rockmass quality, induced stress, structure, or stope orientation could affect stope stability.  This paper presents the fundamental principles of the Stability Graph method, the Python script and workflow, and two case studies to demonstrate the utility of the tool.

Oct 1, 2024

By Glen Frank, David R. Chapman

Determining the number of boulders and cobbles that will be encountered in the tunnel envelope is one of the most difficult problems confronting the geotechnical engineer in the design of tunnels in glacial terrain. This issue is very important to the contractor, both in determining what methods will be used to construct the facilities and the advance rate that can be achieved for the chosen construction methods. The cost of constructing tunnels by pressurized face tunneling techniques is very sensitive to accurate characterization of the very coarse fraction (large cobble and boulder sized clasts), and no standard method is current available for characterizing boulders and cobbles. Careful use of large diameter borings and the sampling and analytical techniques outlined in this paper can provide a means to answer this difficult question. By performing large diameter borings and carefully recording and sampling the cobbles and boulders encountered, an estimate for the amount of these materials present in the formations sampled can be provided. Using this method allows the designer to communicate the expected geologic conditions to the prospective bidders in a manner that is both definitive and defensible. This paper describes the method as it was used in conjunction with the design of a large diameter sewer tunnel in the Columbus, Ohio area.

Jan 1, 2005

By J. B. Mackintosh

The general method which has been followed since the time of Heinrich Rose, and perhaps before, for the determination of phosphorus in iron and steel, is to dissolve the sample either in nitric acid, aqua regia, or some other oxidizing agent, in order to convert the phosphorus into phosphoric acid. In 1832 Rose* recommends HNO3, or aqua regia, preferably the former. Wohler, in 1854,† directs that the iron should be oxidized by burning in a stream of oxygen for the determination of carbon, and the resulting oxide of iron containing all the phosphorus as phosphoric acid be treated as an ore. In the fourth English edition of Fresenius' Quantitative Analysis,† pp. 666-667, we find directions to dissolve the iron in aqua regia or bromine; and, coming down to more recent times, we find F. A. Cairns § directing the employment of a boiling solution of potassium chlorate and hydrochloric acid for the oxidation and solution of the sample to be analyzed. It is unnecessary to multiply quotations from the literature of the subject, though they might be extended almost inclefinitely, since it would only be a repetition of the above mentioned series of methods in one form or another; and the references given have been selected to show how little essential variation there is in the type-methods of solution used. In fact it seems to be universally accepted as a necessary first step in the determination that some such method of solution should be followed. One chemist, however, struck out on a different line of thought, and to the Lest of my knowledge only one, M. V. Tantin, who, in

Jan 1, 1886

By Garth D. Kirkham, James P. Siddorn, Hendri Falck

"Abstract - As an integral part of the Geological Survey of Canada’s EXTECH III Project, a threedimensional computer model has been created of the Con and Giant gold mines in the Yellowknife Greenstone Belt. The Yellowknife Mining Camp has produced a great deal of gold (over 15 million ounces, or 400 000 kg) in the past sixty years and the prevailing assumption is that the mines are at the end of their lives. However, the potential for discovering additional resources using new ideas and technology, and maximizing use of existing data, offers an opportunity to sustain the old mining operations and maintain their contribution to the community.The objectives of this study included the compilation and creation of a three-dimensional Geographic Information System model (combining 3D subsurface data with surface information) to provide a framework for concurrent structural and alteration studies, and to encourage new exploration in the area. A secondary objective was to retain the data in digital format as a historical reference for future remediation efforts, and to document a portion of Canada’s rich mining history.The end result is a database and set of models that have been instrumental in facilitating further exploration in the area, as well as in developing targets for future exploration and mining activities. In addition, the EXTECH III 3D GIS model has made a valuable contribution toward the understanding of the arsenic remediation issues at the Giant mine, and has served as a tool in support of a variety of highly technical engineering studies and for communicating information and problems to the layman and community as a whole."

Jan 1, 2003